Treatment of metal sulphide concentrates by roasting and electrically stabilized open-arc furnace smelt reduction

ABSTRACT

A process for treating a metal sulphide concentrate which includes the steps of: a) roasting the concentrate to reduce the sulphide content of the concentrate, to a negligible value and b) melting the concentrate, under reducing conditions, in an electrically stabilized open-arc furnace, in particular a DC arc furnace.

This is a continuation application of International ApplicationPCT/ZA00/00032 filed Feb. 25, 2000, which claims priority to SouthAfrica 99/1285 filed Feb. 26, 1999.

BACKGROUND OF THE INVENTION

This invention relates to the treatment of metal sulphide concentrates.

PRIOR ART Roasting for Sulphur Removal

Oxidative roasting of pyrite (FeS₂) is a standard way of producingsulphuric acid. Roasting is used on an industrial scale, e.g. for theproduction of zinc, copper, and nickel, even tin, molybdenum, andantimony and, in many cases, takes place in conjunction with one or moreleaching or smelting operations. Sulphide roasting is used to oxidizesome (or all) of the sulphur. The resulting SO₂ is treated further, mostcommonly producing sulphuric acid. Other options for recovery of sulphurinclude the production of elemental sulphur, or liquid SO₂.

Modem roasting processes usually use fluidized-bed reactors, which areenergy-efficient, and have a high productivity because of theirfavourable kinetic reaction conditions. The SO₂ content in the off-gasis typically 8 to 15% by volume.

For pyrometallurgical processing, the usual purpose of roasting is todecrease the sulphur content to an optimum level for smelting to amatte. Partial (oxidizing) roasting is accomplished by controlling theaccess of air to the concentrate; a predetermined amount of sulphur isremoved and, for example in the recovery of copper, only part of theiron sulphide is oxidized, leaving the copper sulphide (for example)relatively unchanged. Total, or dead, roasting involves the completeoxidation of all sulphides, usually for a subsequent reduction process.

There are many modem pyrometallurgical processes in which roasting isnot a separate step, but is combined with matte smelting. Flash furnacesemploy sulphide concentrate burners that both oxidize and melt the feed,and are used extensively in the copper industry. Autogenous bathsmelting is another alternative. Here a lance blows air or oxygen,together with concentrates and reductant, into a molten bath, and theenergy released by the oxidation of the sulphur provides much of therequired energy for the smelting process.

The roasting process has several effects:

a) Drying the concentrates

b) Oxidizing a part of the iron present

c) Decreasing the sulphur content by oxidation

d) Partially removing volatile impurities, for example arsenic

e) Preheating the calcined feed with added fluxes (for example, silicaor limestone), in order to lower the energy requirement of thedownstream process

Environmental concerns have highlighted the need to lower the emissionsof sulphur from smelters treating sulphidic raw materials. Theseemissions emanate primarily from the furnaces and converters, either asfugitive emissions or as process gases vented up a stack. It should benoted that the typical 1 to 2% SO₂ in the off-gas from reverberatoryfurnaces (for example) is too low for effective acid production.

The general trend in recent years has been to eliminate as much aspossible of the iron sulphides (usually pyrrhotite) during the millingand flotation stages, in order to minimize the sulphur input tosmelters.

Dead roasting, i.e. close to 100% sulphur removal, has the benefit ofremoving essentially all the sulphur at the beginning of a smeltingprocess. Furthermore, in comparison with the intermittent nature of SO₂produced in a converting operation, a steady and almost optimum SO₂content of off-gas from a roaster requires a smaller and less expensiveacid plant.

Copper

Various roasting techniques in the recovery of copper are described inthe literature. ¹⁻¹⁶

Copper—Brixlegg

In the Brixlegg process, copper was produced by electric smelting ofdead-roasted chalcopyrite concentrate in a circular AC (alternatingcurrent) submerged-arc furnace, using coal as a reductant.

Brixlegg reports a 95% recovery of copper to blister, and levels ofcopper in the slag of less than 1% have been claimed. The crude copperaveraged only 95% copper, and the operation has been discontinued¹.Disadvantages of this process are the relatively high copper losses inslag, and the high electrical energy consumption.

An undesirable aspect of the Brixlegg process is the fact that leadpasses into the final copper anodes and makes them fragile if theconcentration is too high. On the other hand, the exceptionally highrecovery of other metals related to copper makes the process ofparticular interest for treating ores which contain nickel and noblemetals. (The nickel can be separated from the anode mud.)

A submerged-arc furnace has been used for treating dead-roasted calcinein a process developed by the US Bureau of Mines¹⁴, as was also used inthe Brixlegg process. It was found that in order to produce ahigh-purity blister (2.2% total impurities) and low-copper-content slagin a submerged-arc furnace, a two-cycle procedure was required. Usingthis rather inconvenient and non-continuous procedure, recoveries ashigh as 98% were attained.

Nickel

In the nickel industry, Falconbridge¹⁷⁻²⁴ and Inco²⁵⁻²⁹ have worked onprocesses involving the smelting of roasted sulphide concentrates. Theseprocesses use six-in-line furnaces, commonly employed in that industry,which generally operate at temperatures around 1400° C. The reductionreactions needed to provide the appropriate conditions for recoveringmetals from the oxides tend to raise the operating temperature of thesefurnaces. Consequently, large volumes of air are drawn into the furnaceto cool the freeboard space of the furnace. This tends to result in highlosses of the feed materials as dust. Dust losses of up to 25% of thefeed have been mentioned²⁰.

Nickel production has however been accompanied by a level of SO₂generation which is environmentally unacceptable. It has been recognisedthat a major means to reduce SO₂ emissions is to increase the degree ofsulphur elimination in the fluidized-bed roasters. However, the existingfurnace technology is limited in the degree to which highly roastedconcentrates can be handled. The higher degree of roast demands morestrongly reducing conditions in the furnace to smelt more oxidizedcalcine feed, and to counteract slag losses. Higher coke addition ratesare needed. Extra energy is generated by the additional coke combustionproducts, resulting in a higher temperature in the furnace freeboard.This requires greater amounts of cooling air to control the temperature.The furnace off-gas handling system capacity would have to be expandedto handle the greater quantities of gas. Also, the more metallized mattemelts at higher temperatures, demanding superheated slags to controlmatte temperatures and bottom build-up. Refractory erosion in the slagzone with higher temperature slags must be controlled by cooling therefractory with copper coolers.

About 25% of the calcine escapes the six-in-line furnace; as much aspossible of this is recycled back to the furnace²⁰.

Inco's roast-reduction smelting process²⁵⁻²⁹ involves deep roasting ofnickel concentrate in fluidized-bed roasters. The roaster off-gas istreated in a sulphuric acid plant. The low-sulphur calcine is reductionsmelted with coke in an electric furnace to yield a sulphur-deficientmatte. This sulphur-deficient matte is converted to Bessemer matte inPeirce-Smith converters, with minimal evolution of sulphur dioxide(because of its sulphur-deficient nature), and the converter slag isreturned to the electric furnace. Excellent recoveries of nickel wereobtained, and the process was developed up to commercial-scale testingat the Thompson smelter during 1981 to 1982. Flash smelting of bulkcopper-nickel concentrates was considered superior at Inco's CopperCliff smelter, but it was seen that in other circumstances theroast-reduction process could be an attractive option.

Sulphur is eliminated from the concentrate mainly in the roasters,running at 830 to 850° C. The high temperatures promoted high oxygenefficiency, of approximately 95%. Slurry feeding permitted excellentcontrol of the air to concentrate ratio in the roaster, and good controlof sulphur elimination (approximately 80%). The process resulted inhigher furnace temperatures, as well as higher iron levels to beoxidized in the converters.

U.S. Pat. No. 4,344,792²⁵ describes the possibility of smelting either apartially roasted concentrate or a blend of dead-roasted concentrate andgreen concentrate, together with a carbonaceous reductant and silicaflux. The feed is to contain only sufficient sulphur to produce a matte,in which the iron is present as metallic iron, and which has a sulphurdeficiency of up to 25% with respect to the stoichiometric base metalsulphides Ni3S₂, Cu₂S, and Co₉S₈. The iron is later converted, toproduce a low-iron matte by blowing and slagging the iron with silicaflux, with very little release of sulphur dioxide during this stage ofthe process.

A process for the treatment of pyrrhotite, based on roasting toeliminate all or part of the sulphur, and hydrometallurgical treatmentof the calcine to recover nickel, is described in Kerfoot³⁰.

Platinum Group Metals

Sulphide ore concentrates containing platinum group metals (PGMs) havebeen roasted for various leaching processes.

The US Bureau of Mines devised a procedure for selectively extractingPGMs and gold from Stillwater Complex flotation concentrate. Theconcentrate was roasted at 1050° C. to convert base-metal sulphides tooxides, and the PGMs from sulphide minerals to their elemental states.The roasted concentrate was then treated in a two-stage leachingprocess. Up to 97% of the platinum, 92% of the palladium, and 99% of thegold were extracted from the roasted concentrate³¹.

Other techniques are described in References 32 and 33.

Zinc

Dead roasting of zinc concentrates is practised at industrial scale atZincor, in Springs, South Africa. The calcine from this operation istreated by leaching and electrowinning.

A sulphide concentrate comprising 15% copper, 17% zinc, and 10% lead wasroasted in a laboratory-scale fluidized bed in China, with the intentionof using the product for further hydrometallurgical or direct smeltingprocessing³⁴.

Prime Western grade zinc has been produced from lead blast-furnace slags(and other zinc-containing waste materials) at large pilot-plant scaleat Mintek in Randburg, South Africa, using the Enviroplas process³⁵.Feed materials are smelted in a DC arc furnace, and the zinc is fumedoff as a vapour, leaving behind a slag containing only small quantitiesof zinc oxide. The zinc vapour is subsequently treated in a lead splashcondenser, resulting in the production of Prime Western grade zinc.

SUMMARY OF INVENTION

The invention provides a process for treating a metal sulphideconcentrate which includes the steps of:

a) roasting the concentrate to reduce the sulphide content of theconcentrate, and

b) smelting the concentrate, under reducing conditions, in anelectrically stabilized open-arc furnace.

As used herein the phrase “an electrically stabilized open-arc furnace”means a DC arc furnace or an electrically stabilized single electrodeopen-arc AC furnace (SOA furnace).

Preferably the roasting reduces the sulphide content to less than 10%sulphur by mass and, more preferably, to less than 1% of the initialamount present, the objective being to reduce the sulphide content to anegligible or otherwise acceptable value. The material produced by thisstep is referred to herein as “highly-roasted” or “dead-roasted”.

Preferably the roasting is done in a way which produces a steady streamof SO₂-bearing gas. This gas may be used as a feedstock in a sulphuricacid plant. This step may be implemented in any appropriate way and forexample the roasting may be performed in an enclosed vessel such as afluidized bed reactor, to provide a high-concentration of SO₂ in thegas.

Alternatively the SO₂-bearing gas which is released in the roastingprocess may be subjected to gas scrubbing and neutralization.

The elimination of sulphur results in the valuable metals beingcollected in an alloy or from a vapour rather than as a matte, duringthe following smelting stage. This is believed to be advantageous asalloys have a greater collection efficiency than mattes.

The term “alloy” is used here to denote a mixture of metals which may ormay not contain some sulphur, as distinct from “matte” which is amixture of metal sulphide.

The aforementioned process may be varied according to requirement and,more particularly, according to the nature of the metal sulphide ormetal sulphide which are being treated.

When used for the treatment of zinc sulphide the calcine from thefluidized bed reactor may, optionally, be agglomerated before being fedto the arc furnace.

The reductant used in the furnace may be of any suitable kind and mayfor example be coke.

Zinc in the calcine is reduced to metal vapour and may be fumed off in agas stream for recovery in any suitable way, for example in a leadsplash condenser.

The aforementioned process as used for the treatment of zinc sulphide isparticularly suitable for the treatment of zinc concentrates and zincore which contain relatively high levels of manganese, for example asencountered in the Gamsberg deposit in South Africa which has amanganese content which is higher by a factor of about 10 than themanganese content normally encountered in zinc concentrates.

The process may also be used for the treatment of nickel, copper andcobalt sulphide concentrates, whether existing as separate or combinedsulphide, and PGM concentrates. The PGM concentrates may also be in theform of green furnace matte which is produced in any appropriate way,for example by making use of a six-in-line electric furnace.

The smelting of the dead-roasted concentrate, in the arc furnace,produces a slag which is depleted in metal values and an alloy. Iron inthe alloy may optionally be removed, in oxide form, from the alloy usingany suitable technique, such as by making use of a Peirce-Smithconverter or a Top Blown Rotary Converter (TBRC).

Other undesirable elements such as carbon, silicon, or chromium, may beremoved from the alloy using any suitable technique such as converting,or refining in a ladle furnace for example.

If the converter is used then alloy from the converter or, otherwise,alloy drawn directly from the arc furnace, if the converter is not used,is atomized so that it is in a form which is suitable for subsequenthydrometallurgical recovery of metal values, e.g. using a suitableleaching process.

Atomization of the alloy solves the problem of having to crush and millan extremely tough alloy.

The green furnace matte may be granulated and milled, or water atomized,prior to the roasting step.

The smelting step may be a two-stage reduction smelting process,particularly when treating concentrates containing appreciablequantities of nickel and copper.

In the first stage, use could optionally be made of an arc furnace whichis operated under slightly reducing conditions. This stage allows forthe settling of some of the copper and nickel in an alloy, and a largefraction of the PGMs which partitions to the alloy.

In the second stage, which may be carried out in an arc furnace which isoperated under highly reducing conditions, substantially all of theremaining nickel, the remaining PGMs, and most of the cobalt, areremoved in an iron-based alloy which may also contain some copper.

The iron-based alloy may be atomized in preparation forhydrometallurgical treatment, e.g. for treatment in a leaching step.

The copper/nickel alloy from the first stage may be prepared forhydrometallurgical recovery by being water or gas atomized, granulated,or crushed and milled.

In carrying out the process of the invention use is made of a stabilisedopen arc furnace. A suitable furnace of this type is a DC arc furnace.The invention is however not limited in this regard for it may bepossible to stabilise the arc or arcs of an AC arc furnace, usingsuitable control techniques, to achieve characteristics which aresimilar to those of a DC arc furnace in that the arc, or each arc,extends vertically from an overhead electrode to the charge, isconfined, and does not deflect to side walls of the furnace.

Smelting in a DC Arc Furnace

A stabilized open arc furnace offers a number of advantages in thesmelting of roasted sulphide concentrates and is seen as the enablingtechnology to make possible the process of the invention.

A DC arc furnace is roughly cylindrical in shape, often having a conicalroof. A single vertical graphite electrode is used as the cathode, andthe anode is embedded in the bottom of the furnace, in contact with themolten bath. The usual configuration involves operation with an opentransferred plasma-arc above a molten bath with a surface substantiallyuncovered by feed materials (ie. an “open bath” operation). However,work has also been done using a two-electrode configuration (atwin-cathode is sometimes used for steel scrap or DRI (direct reducediron) melting, and a two-electrode cathode-anode arrangement has alsobeen used on a pilot scale.) Feed materials are either fed through thecenter of the electrode, or through a feed port fairly close to theelectrode. Fewer feed ports are required with this configuration offurnace than are normally required for an AC six-in-line or a threephase three electrode AC furnace.

The powerful concentrated plasma arc jet provides a very efficient formof energy transfer to the molten bath of the furnace. This enablesreactions to take place fairly rapidly, and good mixing is establishedin the bath, leading to a fairly uniform temperature distribution. TheDC arc is relatively stable, not too easily extinguished, and isdirected downwards towards the molten bath, with little flare towardsthe furnace walls. The arc jet ‘pulls’, via the Maecker effect, thefurnace gases towards it, thereby attracting fine feed materialsdownwards into the bath, in so doing minimizing dust losses from thefurnace. The low gas volumes from an electric furnace (compared to afurnace where energy is provided by combustion) also help in minimizingdust losses. The DC arc furnace can handle fine feed materials,typically sized below 3 mm, which makes it well suited for coupling to afluidized-bed roaster.

The simple configuration of the DC arc furnace allows the freeboard tobe well sealed, maintaining the CO atmosphere intemally, and minimizingthe ingress of air.

Very high operating temperatures (much higher than those usuallyencountered in conventional base metals smelting) can be attained in thefurnace, if required by the process, as power is supplied by the openarc, not merely by resistance heating in the slag. The furnace roof andwalls are cooled (for example, by water-cooled copper panels) to retainthe integrity of the furnace, even under conditions of high-intenisitysmelting. Sigh freeboard temperatures are easily accommodated. Thepossibility of strongly reducing conditions in the furnace (togetherwith the high operating temperature) avoids the common difficulties withthe build-up of high-melting magnetite leading to operational problemsin the furnace.

The processes described herein have high recoveries of the desiredmetals, and produce very clean slags. The DC arc furnace works wellusing iron alloy collection of valuable metals, or fuming off volatilemetals. The processes result in low levels of impurities in the desiredproducts.

The application of a DC or SOA arc furnace provides unique advantagesparticularly for feeds that contain high amounts of iron oxide whichrequires lots of reduction and for feeds that contain, for example,nickel and cobalt which require low oxygen potentials to achieve lowslag losses.

A comparison of the characteristics of conventional furnaces andstabilised open arc furnaces highlights the advantages of usingstabilised open arc furnaces in the process of the invention.

Conventional Furnaces

A conventional furnace has limitations in handling CO gas in thefreeboard. Sealing the furnace is very difficult with multipleelectrodes and feed points and a large cavity for the off-gas system.Rather than attempting to seal the furnace, the standard design involvesthe addition of air to combust the CO in the furnace freeboard and theaddition of even more air to temper the combustion product gases. Thisresults in large off-gas volumes, large quantities of dust make and theneed to operate fans in a dirty gas environment.

In reduction smelting, reductants such as coke are mixed into thecalcine. The reduction reaction is relatively slow in a conventionalsmelting configuration where the calcined material is smelted on a slagbath surface. The power density of the furnace which corresponds to thesmelting rate cannot exceed the reduction reaction rate. This limitationbecomes more important as the degree of roast increases and as theamount of reduction increases.

A calcine bed resting on the slag, or material banked up on the sidewalls, can roll over into the slag bath and cause unwanted slag foaming.

A conventional furnace (a rectangular six-in-line furnace, for example)requires good distribution of feed over the surface of the slag bath.This requires numerous feed points and a complex feed system above thefurnace.

The reduction reaction is dependent on the reductant type. Generally,the need is for fine coke to maximize the reduction reaction rate.Otherwise, coke accumulates at the slag-calcine interface and redirectsthe furnace power, undesirably.

The metallized matte or metal that results from reduction smelting has ahigher liquidus temperature. This necessitates a higher matte or metaltemperature. A conventional furnace has poor capability to transferenergy in the vertical direction between the slag and matte phases.

Stabilised Open Arc Furnace

A DC or SOA arc furnace:

(a) is small and intense;

(b) has no obvious limit on coke reduction kinetics. The ultimate caseof dead roasting and back reduction to alloy is easily possible;

(c) can use a wide range of reductants eg. coal or coke of various sizeranges;

(d) is easy to seal to contain a CO atmosphere, has little offgas,little dust, and few feed points; and

(e) produces hot metal

Recent developments in electrical power supply equipment have resultedin the possibility of using a three phase AC system to provideelectrical energy via a single (graphite) electrode. This can beachieved by the switching of the three phase supply, possibly usingpulse width modulation techniques, to generate a high frequencysynchronized AC output. This development implies that a furnaceconfigured in a similar manner to a DC arc furnace can be designed andused for similar process applications. Robicon of the USA have a powersystem (The Harmony Series) that can provide AC power as described aboveas well as the usual DC power output.

The stability of the AC high frequency arc is claimed to be better thana DC arc. There are a few possible disadvantages of the AC system:

the graphite electrode current capability may be less (skin effect);

the electrode wear may well be somewhat higher than with a DC furnace;

the arc jet and hence heat transfer to the bath may be slightly lowerthan for a DC arc furnace; and

the high frequency may generate harmonics although with suitable solidstate switching techniques the harmonics may be reduced.

Potential benefits of an AC single electrode high frequency arc furnaceinclude:

an improved arc stability (the arc is less likely to extinguish undercertain circumstances); and

the arc may be less susceptible to loss of vertical directionality (e.g.sideways deflection) due to magnetic effects (deflection of the arc candamage the furnace sidewall and increase the energy losses).

Thus a suitably controlled power system can generate a high frequencywaveform derived from a 3 phase alternating power supply which can bedirectly impressed across a single graphite electrode and a charge in afurnace, to produce a single open arc which is analogous to an arc in aDC furnace. This arrangement can offer similar and, in some regardsimproved, characteristics compared to a DC furnace and the scope of theinvention therefore extends to the use of a DC arc furnace or astabilised open arc AC furnace in the reduction smelting step.

BRIEF DESCRIPTION OF THE DRAWINGS

The invention is further described by way of examples with reference tothe accompanying drawings in which:

FIG. 1 is a generalised flow chart of process steps according to theinvention for the treatment of various metal sulphide concentrates,

FIG. 2 illustrates a variation of the generalised process of FIG. 1suitable for the treatment of zinc concentrates,

FIG. 3 illustrates a variation of the generalised process of FIG. 1suitable for the treatment of PGM and nickel and copper sulphideconcentrates,

FIG. 4 is a general flowsheet of high roast-reduction,

FIG. 5 is a particular example of the process of FIG. 1 for thetreatment of PGM matte concentrates, and

FIGS. 6 to 10 depict hydrometallurgical treatment procedures which canbe used in the process of the invention, the choice of a particularprocedure depending on the elements of interest in the alloys which areproduced in a preceding pyrometallurgical phase of the process.

DESCRIPTION OF PREFERRED EMBODIMENTS

Zinc concentrates of the kind encountered in the Gamsberg deposit have amanganese level which is up to 10 times higher than normal. This highmanganese level causes problems and additional costs when recovering thezinc, after leaching, in a conventional electrowinning plant For theelectrowinning route much research has been carried out on means ofremoving the manganese from the electrolyte, or on electrolyticprocesses which enhance the production of MnO₂ at the anode in a zinccell. The former technique is expensive, and the latter approach, whichis directed to the production of high quality electrolytic manganeseoxide, appears to be problematical.

It is advantageous to remove sulphur from metal sulphide concentratesbefore smelting. For example existing PGM and base metalpyrometallurgical processes have a number of limitations, particularlyin the converting stage. It is difficult to achieve environmentallyacceptable levels of sulphur capture, especially in view of the problemof fugitive emissions from Feirce-Smith converters. Converting is abatch process which has inherent scheduling problems, losses andspillages from the crane transport of ladles, and high labour costs.

Another consideration with conventional furnaces is that there is alimit to the amount of PGM-containing UG2 concentrate that can betreated. A problem with chromite in UG2 concentrate is that it cannoteasily be solubilised in slag during normal smelting. A spinel forms,builds up in the furnace, and needs to be dug out frequently.

The invention is described hereinafter firstly with reference to ageneralized treatment process, as exemplified in FIG. 1, and thereafterwith reference to three particular forms of the process shownrespectively in FIGS. 2, 3 and 5.

FIG. 1 of the accompanying drawings illustrates a generalized processfor the treatment of metal sulphide wherein a concentrate 10 of themetal sulphide or metals sulphide is fed to a fluidized bed roaster 12which, preferably, produces a steady stream of high strength SO₂ bearinggas 14 which can be used as feedstock for example in a sulphuric acidplant. This is not essential to the process though, for gas from thereactor could alternatively be subjected to gas scrubbing andneutralization.

The calcined product from the roaster is fed to a DC arc furnace 16together with a reductant 18 which, for example, is in a form of coke.

In the fluidized bed reactor 12 the sulphur content of the concentrateis reduced substantially, to approximately 10% by mass in the case ofhigh roasting, or to approximately less than 1% of the initial value, oreven lower, in the case of dead-roasting. The elimination of sulphurfrom the concentrate results in the valuable metals being collected inan alloy 20 which is produced by the furnace 16, rather than as a matte.Alloys have a much greater PGM collection efficiency than matte.

The furnace also produces a slag 22 which is depleted in metal values.

The nature of the process thereafter depends on the nature of theconcentrate which is being treated. If the concentrate is a zinc-bearingconcentrate such as zinc sulphide then the zinc is reduced to metal inthe DC arc furnace and fumed off in a gas stream or vapour 24 which ismainly zinc and carbon monoxide. The gases are led directly to alead-splash condenser 26 for absorption, or condensing, and subsequentrecovery as a product 28.

For the treatment of copper, nickel or cobalt sulphide, or PGM sulphide,the alloy produced by the DC arc furnace may be atomized (step 30) andthen subjected to a hydrometallurgical recovery process 32 to produce aproduct 34. Alternatively the alloy is first fed to a converter 36 suchas a Peirce-Smith converter and then to the atomizer 30 and through tothe hydrometallurgical process 32.

The converter slag 37 may be returned to the DC arc furnace 16.

The process 32 may be of any appropriate type and a particularlysuitable process 32, which is intended to fall within the scope of theinvention, is described hereinafter with reference to FIG. 5.

FIG. 2 is a particular example of the process of FIG. 1 for the recoveryof zinc from a high manganese ore such as for example the Gamsberglead-zinc deposit.

It is assumed that the mining of ore from the Gamsberg deposit, followedby grinding and flotation, yields a concentrate 10 which contains about48% zinc, 29% sulphur, from 4% to 5% manganese as oxide, and 5%moisture. The concentrate 10 is fed from a suitable store to fluidizedbed roasters 12 where the sulphur content of the concentrate is reducedto approximately 0,75%. The gases from the roasters are cooled in awaste heat boiler, cleaned in cyclones, subjected to electrostaticprecipitation, and are then passed to a sulphuric acid plant 14. It isassumed that the final exhaust gases can be discharged to atmospherewithout the need for scrubbing out the last traces of sulphur dioxide.

The calcine which contains about 58% zinc as oxide is fed with dry coke18 and a small amount of lime to the DC arc furnace 16, with a sealedfreeboard. The calcine may optionally be agglomerated in a step 38before being fed to the DC arc furnace. This step does however involveadditional capital and operating costs.

In the DC arc furnace the zinc oxide is reduced to metal and fumed in agas stream 40 which principally contains zinc and carbon monoxide. Thesegases are led directly to the lead splash condenser 26 where the zincand any lead are removed from the gas stream by absorbing or condensingthese metals in a curtain of lead droplets. The gases 42 exiting thecondenser are burnt in a combustion chamber, cooled in a waste heatboiler and are cleaned in a bag filter 44 before being exhausted toatmosphere. The maximum concentration of sulphur dioxide in the exhaustgases is estimated to be less than 100 parts per million which does notpose an environmental problem. The dust collected in the bag filter,which consists mainly of zinc oxide, is washed with water to remove anyhalides before being returned to the roasters.

Slag 46 which is produced by the furnace is granulated before beingremoved to a waste storage dump. A small amount of metal 48 produced inthe furnace is periodically tapped from the furnace and is run intorough moulds.

Dross 50 from the lead splash condenser 26 is collected and batchtreated in a small furnace to separate out lead. The dry dross is thenrecycled to the DC arc furnace. Impure zinc 52 from the lead splashcondenser is transferred by ladle to a zinc distillation plant holdingfurnace. A plant of this type requires a constant feed and consequentlyprovision is made for casting ingots or reheating ingots to balance theflow from the condenser so that the requirements of a zinc distillationplant 54 are met. The distillation plant produces SHG (Special HighGrade) zinc 56, an impure lead and cadmium-zinc alloy 58 which istreated according to requirement, and a small amount of hard zinc 60, inthe form of a Zn—Pb—Fe—Cu alloy, which is recycled to the DC arcfurnace.

FIG. 3 illustrates the steps of a particular form of the process shownin FIG. 1 used for the treatment of PGM and base metal concentrates 10.

The feed material 10 is dead-roasted in a fluidized bed roaster 12 whicheffectively reduces the sulphur content of the concentrate to zero. Thislimits later sulphur emissions from the concentrate. The reactor 12produces a steady stream of SO₂-bearing gas which is used as feedstockfor a sulphuric acid plant 14. It is to be noted though that theSO₂-bearing gas may alternatively be subjected to gas scrubbing andneutralization instead of being used as feedstock.

The concentrate, after being roasted, is fed to a DC arc furnace 16 toproduce an alloy 20, and a slag 22 which is depleted in metal values andwhich is discardable.

It may possibly be beneficial to add a base metal collector 62, as isindicated in FIGS. 1 and 3 in dotted outline, to the fluidized bedreactor 12. For example by adding nickel (e.g. in the form of nickelsulphate) or copper (e.g. in the form of copper sulphate) to thefluidized-bed roaster, along with the concentrate, a greater quantity ofnickel or copper (as oxide) is established in the furnace feed, and thisdecreases the requirement to reduce a large quantity of iron which wouldotherwise be required to produce sufficient alloy for effectivecollection of the valuable metals.

Any base metal oxide, sulphate or sulphide, which is compatible with theprocess and which is in a fine form which can react with the feed, couldbe used as a collector.

The alloy 20 may directly be passed to an atomizer 64 which makes thealloy suitable for subsequent leaching in a hydrometallurgical step 32,for the recovery of metal values. Alternatively the alloy is fed to aconverter 36 which removes most of the iron from the alloy in oxideform. The resulting slag 37 may be returned to the DC arc furnace 16.Thereafter the alloy is atomized and subjected to the aforementionedleaching step to enable the metal values to be recovered.

Disadvantages or problems which are overcome or reduced are thedifficulty of hot ladle transportation from the furnaces to theconverters which create scheduling problems in the converter aisle,losses due to spillages, skull formation in the ladles, high labour andmaintenance costs as well as pollution problems. Advantages include theelimination or reduction of the Cr-spinel problem in the furnace, thetolerance for higher Cr-levels in the feed, with resulting higher PGMrecoveries, the avoidance of the matte breakout problem, a lowering ofenergy consumption and, in one variation of the process, the eliminationof the converter.

The process is also able to accommodate a wide range of feed materials,up to 100% UG2 (in the case of PGM concentrates), with a higher chromitecontent. This provides significant advantages with respect to PGMrecovery in the mining and concentrating operations.

FIG. 4 illustrates a general flowsheet of high roast-reduction smeltingof base-metal sulphide concentrate which incorporates a DC arc furnace.It is understood that there are a number of variations of thisflowsheet.

Concentrate slurry 66 is continuously fed to fluid bed roasters 68. Thedegree of concentrate sulphur elimination (degree of roast) may varyfrom 70% all the way to 100% (i.e. dead-roast). Roaster off-gas 70 iscooled, cleaned, and directed to an acid plant for SO₂ fixation. Calcine72 premixed with flux and coal 74 is smelted in a DC arc furnace 76. Thesmelting furnace would produce discard quality slag 78. The grade, ironand sulphur content of the alloy or matte 80 produced will depend on thedegree of roast and the ratio of reductant to calcine. The furnaceoff-gas 82 would be of a low volume and high CO concentration. A highgrade, low sulphur, highly metallized matte from the furnace wouldrequire minimal treatment in a pyrometallurgical converting process 84.

The following is a list of variations of the general flowsheet in FIG.4:

pre-reduction (solid state) of the calcine in an external vessel;

pre-heating of the calcine in an external vessel;

reuse of the DC furnace off-gas for either of the above;

direct hydrometallurgical refining of the DC furnace alloy/matte; and

cleaning of the converter slag in a suitable slag cleaning process.

The DC arc furnace can easily accommodate the high degree of oxidereduction required. The DC arc furnace also allows for the production ofmore alloy as more reductant is added to the furnace.

FIG. 5 illustrates a particular form of the technique of FIG. 1 wherein,in a continuous process, sulphur is removed by the dead-roasting ofPGM-containing matte followed by a smelting stage under reducingconditions.

A six-in-line or a three-electrode furnace is used for the production ofPGM-containing matte from concentrate. The green furnace matte is thengranulated and milled, or water atomized, to produce a feed 10 which isfed to the fluidized bed roaster 12. A steady stream of gas containingSO₂ is fed to a sulphuric acid plant 14.

The roasted material is then subjected to a two-stage reduction smeltingprocess which makes use of a first furnace 86 and a second furnace 16,which is a DC arc furnace. The first furnace may be of any appropriatetype and may for example be a DC arc furnace. The first furnace allowsfor the settling, under slightly reducing conditions, of some of thecopper and nickel as an alloy 88. A large fraction of the PGMspartitions to this alloy which is then treated in an atomizer 90 beforebeing directed to a hydrometallurgical process 32.

Slag 92 from the furnace 86, and a reductant 18 are fed to the secondfurnace 16 which operates under highly reducing conditions in order toremove virtually all of the nickel and PGMs contained in the slag, alongwith most of the cobalt, to produce an iron-based alloy 94 which mayalso contain some copper. This alloy passes to a converter 95 and isthen atomized in a step 96 in preparation for leaching in ahydrometallurgical process 32. Slag from the converter may be returnedto the furnace 16.

Slag 98 produced by the DC arc furnace is discardable and issufficiently devoid of valuable metals that it can either be discardedor used in applications such as road construction or shot blasting.

Again it should be noted that a base metal collector 62 could be used,as has been described hereinbefore, to decrease the capability whichwould otherwise be called for, of the furnace to reduce a large quantityof iron.

A simpler and potentially more cost effective process to theaforementioned two-furnace process involves the single-stage smelting ofthe roasted furnace matte in a DC arc furnace. This is essentially thetechnique which is shown in FIG. 3, where the feed material 10 is agreen furnace matte. The simpler process requires fewer process unitsbut it has the disadvantage of capturing all base metals and some irontogether with the PGMs.

A number of examples of the invention have been described hereinbefore.In each case the concentrate, which optionally is in green furnace matteform, is dead-roasted and thereafter is smelted under reducingconditions. The essentially complete removal of sulphur means that latersulphur emissions are limited. The spine problem (in the smelting ofhigh chromite containing PGM concentrates) is reduced and discardableslags are produced.

The feed to the DC arc furnace is pre-heated in the fluidized bedreactor. It is to be noted that a DC arc furnace is well suited tohandling fines. The alloy which is produced by the DC arc furnace can bewater atomized (step 64).

Although it is possible to make use of a converter it is also possibleto eliminate the need for a converter and in this way the likelihood ofspillages is reduced and scheduling problems are also reduced.

The nature of the hydrometallurgical process 32 depends on the majorelements of interest in the alloys which are produced by means of any ofthe aforegoing pyrometallurgical techniques. Typically these elementsare iron, nickel, copper, cobalt and PGMs. The hydrometallurgicalprocessing of these alloys depends on case-specific factors. The unitoperations that could be applied in the treatment of the alloy includeambient-pressure leaching, pressure leaching, precipitation, solventextraction, electrowinning and crystallisation. The principles of theindividual unit operations are generally known in the industry. They maybe used in a wide variety of combinations, and a person skilled in thepractise of hydrometallurgy will be able to devise an appropriatecircuit for any specific case.

Each of the Examples shown in FIGS. 6 to 10 embodies a general approach,and is not meant to limit the applicable hydrometallurgical option forprocessing the alloy in question. Process steps for the removal ofimpurity elements such as selenium are omitted for the sake of brevity,but it should be understood that they would be incorporated asnecessary, as known to those skilled in hydrometallurgy. Where acidaddition is shown, it may be either fresh acid or acid recycled from ametal recovery stage such as electrowinning.

The copper solvent extraction and electrowinning stages are asconventionally practised in the industry.

The iron precipitation can be done at elevated pressure and temperature,such that hematite is precipitated and acid is regenerated for recycle.It could also be done by means of neutralisation with an appropriatealkali (an example is limestone, but a number of others exist) such thatgoethite, jarosite, basic ferric sulphate or other similar compound isprecipitated.

The solutions containing cobalt and/or nickel (shown as proceeding toNi/Co separation and recovery) would be treated in the same way as isdone in conventional base metal refining, for the recovery of the cobaltand/or nickel. This could entail the precipitation of cobalt(III)hydroxide or the solvent extraction of cobalt, and the electrowinning ofnickel and/or cobalt. Alternatively, it could entail the crystallisationof mixed or separate cobalt and/or nickel salts, or the precipitation ofhydroxides, sulphide or carbonates. Ion exchange could also be used insome cases.

The examples in FIGS. 1 to 5 have been described with reference to theuse of a DC arc furnace. This is non-limiting for, as has been indicatedhereinbefore, a DC arc furnace is a particular form of a stabilised openarc furnace. Although use of a DC arc furnace is preferred and theoperation of a furnace of this type is well established it is possibleto make use of an AC open arc furnace which has been stabilised, usingsuitable control techniques, to confine the arc in the furnace so thatit extends vertically from an overhead electrode and does not diverge toside walls of the furnace.

EXAMPLE 1 FIG. 6

This Example applies to those situations in which the alloy containsPGMs and valuable base metals. In the first step, the iron and basemetals are dissolved, leaving a residue that comprises a PGM concentratethat can proceed to a PGM refinery. Oxidative leaching would normally beused, but non-oxidative leaching may also be used (in which case theair/oxygen supply to the leach would be omitted). Elevated temperatureand pressure may be used, either alone or in combination withambient-pressure leaching. In some cases, elevated pressure may not benecessary. The resulting solution could be passed directly to a coppersolvent extraction and electrowinning sequence for copper recovery, orit could be passed to an iron-precipitation stage and then to the coppersolvent extraction and electrowinning stage. The raffinate from coppersolvent extraction would be neutralised and any remaining ironprecipitated, to produce a solution containing mainly nickel and/orcobalt, from which these metals can be recovered.

EXAMPLE 2 FIG. 7

This Example applies when PGMs are not present in the alloy, for examplewhen the alloy comes from the reduction of converter slag, for therecovery of base metals. Often, this will entail cobalt as the majormetal value. In this case, the oxidative leach would be operated so asto solubilise the copper, nickel and cobalt while rejecting all or mostof the iron as a hematite or goethite residue. After copper solventextraction and electrowinning, the cobaltinickel solution would proceedto conventional treatment for the recovery of nickel and/or cobalt.

EXAMPLE 3 FIG. 8

In this Example an atomised alloy from a smelting plant is fed to anatmospheric leach, where the bulk of the iron and nickel is leached inthe presence of oxygen and sulphuric acid, at a temperature between 30°C. and 95° C. The copper from the electrowinning spent recycle iscemented in the atmospheric leach and assists in the leaching of theiron and nickel. Conditions for the atmospheric leach were optimisedduring a laboratory scale test programme. A pilot-scale (100 L) batchatmospheric leach, based on the optimised conditions, was performed on5.5 kg of atomised alloy. The performance of the pilot-scale batch leachis summarised below.

Element Feed, g/l Filtrate, g/l Alloy, % Residue, % % Leached Fe 5.6549.60 58.5 4.12 98.9 Ni 105.2 148.4 28.2 4.53 97.6 Co 0.224 0.692 0.580.07 98.2 Cu 2.200 3.660 12.9 71.6 16.4 H₂SO₄ 87.2 1.2 — — —

The leach residence time is set according to the material to limit theleaching of copper while still maintaining high iron and nickelrecoveries. Leach residence times of between 5 and 10 hours arerequired. The optimum leach residence time was exceeded in the testabove, such that some copper leaching was observed.

The residue from the atmospheric leach is then subjected to atwo-stadium pressure leach to remove all the copper and the residualiron and nickel in the presence of sulphuric acid. The pressure leachwas tested in laboratory-scale batch autoclaves. The pressure leachoperates at temperatures between 110° C. and 170° C. with no oxygen inthe first stadium and 0.1 to 6 bar oxygen in the second stadium.Residence times of 60 to 180 minutes are required in the first stadiumand 5 to 60 minutes in the second stadium. The pressure leach residuecontains high levels of PGMs and is suitable for further processing. PGMloss to the leach liquor can be minimised to less than 5% whileachieving a PGM concentrate of greater than 60% precious metals. Thecomposition (mass %) of the PGM concentrate produced from pressureleaching of the atmospheric leach residue is shown below.

PGM + Au Pt Pd Rh Ru Ir Au 61.4 34.6 12.7 4.3 7.7 1.67 0.61 Fe Ni Cu SiCr Se Te S C As 3.6 0.27 3.3 1.75 2.5 0.016 0.007 2.0 0.92 0.49

The composition of the pressure leach liquor is shown below.

Pt Pd Rh Ru Ir Au Fe Ni Cu Ppm ppm ppm ppm ppm ppm g/L g/L g/L <1 16.58.3 7 2.1 <0.5 2 1.8 40

The solution from the atmospheric and pressure leach is treated inpressure vessels to oxidise the iron and precipitate it as hematite.Acid is produced during the hematite precipitation, and the bulk of thesolution following the hematite precipitation from the atmospheric leachliquor is recycled back to the leach. Batch hematite precipitation testswere performed on a laboratory scale to test the removal of iron fromthe atmospheric leach liquor. The pressure oxidation operates attemperatures between 140° C. and 200° C., with oxygen overpressures of 1to 10 bar. The performance of the laboratory-scale batch pressureoxidation is summarised below.

Feed Fe removal Ni Loss [Fe] in Composition, g/L (solid basis) (solidbasis) Filtrate [H₂SO₄] in Fe²⁺ Ni % % g/L Filtrate g/L 24 100 88.9 0.382.6 35.9 35  95 80.5 0.41 6.67 47.4

A bleed stream is taken from the solution following hematiteprecipitation. This bleed is neutralised with lime and any residual ironis precipitated as goethite. The neutral solution is then crystallisedto produce nickel sulphate. The gypsum/goethite residue is disposed of.

The copper sulphate solution from the pressure leach is also treated ina pressure vessel to remove iron as hematite. Selenium is removed in anadditional unit operation to produce a purified solution from whichcopper is electrowon. The spent copper electrolyte is recycled back tothe atmosphere and pressure leaches to utilise the acid generated duringelectrowinning. The copper in the solution is cemented as copper metaland aids in the leaching of the iron and nickel.

EXAMPLE 4 FIG. 9

In this variation the first leach (at ambient and/or elevated pressure)is operated so as to dissolve only nickel and cobalt. The iron andcopper are dissolved and then re-precipitated as goethite and antlerite,respectively. This requires an alloy that is reactive enough to raisethe pH of the leach solution sufficiently for the copper to hydrolyseand precipitate as antlerite. The solution proceeds to nickel/cobaltrecovery. The goethite/antlerite is re-leached to selectively dissolvethe antlerite without co-dissolving more than a small part of thegoethite. The copper-rich solution is passed to copper electrowinning,and the spent electrolyte returned to dissolve more antlerite. Theremaining goethite is then redissolved under more aggressive conditions,leaving the PGMs as a concentrate that is sent to a PGM refinery. Thesolution leaving the goethite dissolution stage is passed to ahigh-temperature autoclave to precipitate the iron as hematite andregenerate acid for recycle to the goethite dissolution stage.

EXAMPLE 5 FIG. 10

This is similar to Example 4, but in this case no PGMs are present,therefore the goethite re-dissolution stage is omitted because it is notneeded.

Test Results

Nickel

Using a DC arc furnace with an internal diameter of 1.0 m, connected toa 5.6 MVA power supply, approximately 26 tons of calcine (‘dead-roasted’concentrate) was processed over a period of 9 days, during which time 83slag taps were carried out. The metallurgical data presented here is aweighted-averaged summary of the operation during 22 taps under thepreferred conditions for producing good metallurgical performance, i.e.just over a quarter of the campaign. These taps cover a wide range ofoperating conditions, but the overall average is consideredrepresentative of the steady operation of the furnace during thiscampaign.

The anthracite addition was approximately 12% based on the mass ofcalcine fed. (Actual additions were 12.7%, 11.6%, and 12.0% during thethree periods summarised here.) Metal was produced at a rate of 250 kgper ton of calcine fed.

Typical operating conditions included feedrates of around 220 kg/h ofcalcine, power levels around 300 kW (including losses of about 150 kW),voltages between 175 and 250 V, and total power fluxes around 400 to 500kW/m². The energy requirement of the process was 760 kWh/t of calcine,excluding losses from the furnace.

Slag Composition During Taps of Good Metallurgical Operation (Mass %)

Fe/ Taps Temp ° C. Al₂O₃ CaO Co Cr₂O₃ Cu FeO MgO Ni SiO₂ SiO₂ 40-46 14816.50 2.85 0.11 1.23 0.46 47.33 8.52 0.26 32.78 1.12 49-56 1483 9.18 2.580.13 1.23 0.51 48.29 6.22 0.33 31.57 1.19 63-69 1505 7.79 2.42 0.13 1.510.48 47.14 5.98 0.29 33.31 1.10 Over- 1489 7.84 2.62 0.12 1.31 0.4947.61 6.94 0.29 32.51 1.14 all

Metal Composition During Taps of Good Metallurgical Operation (Mass %)

Taps Temp ° C. Co Cr Cu Fe Ni S Si 40-46 1470 1.37 0.05 19 33 42 1.30.05 49-56 1450 1.47 0.05 19 32 44 1.3 0.06 63-69 1450 1.60 0.05 19 3342 1.4 0.06

Representative Compositions of Calcine, Slag, and Metal (Mass %)

Calcine Slag Metal Al₂O₃ 3.07 7.84 — CaO 1.43 2.62 — Co (0.46) (0.12)1.5 CoO 0.59 0.16 — Cr — — 0.05 Cr₂O₃ 0.07 1.31 — Cu (5.21) (0.49) 19Cu₂O 5.87 — — CuO (6.52) 0.61 — Fe (32.36) (37.01) 33 Fe₂O₃ 46.27 — —FeO (41.63) 47.61 — MgO 1.78 6.94 — Ni (11.36) (0.29) 43 NiO 14.45 0.37— S 0.78 1.4 Si — — 0.06 SiO₂ 20.38 32.51 — Total 95.4 100.0 98.0Fe/SiO₂ 1.59 1.14 —

Recoveries

The recoveries of the valuable elements were calculated based on thefollowing analyses. The rest of the compositions and flowrates werecalculated on the basis of these numbers.

Typical Best % Co in slag 0.12 0.09 % Cu in slag 0.49 0.43 % Ni in slag0.29 0.20 % Fe in metal 33 33

The actual recoveries obtained on this campaign were calculated usingboth the typical and the best analyses obtained.

Typical Best Co recovery, % 83 87 Cu recovery, % 94 95 Fe recovery, % 2525 Ni recovery, % 98.3 98.9

PGM ConRoast

Approximately 30 tons of PGM-bearing sulphide ore concentrate wastreated in a fluidized-bed reactor, then smelted in a pilot-scale DC arcfurnace. The resulting alloy was refined using a blowing operation, thentreated hydrometallurgically to produce a high-grade PGM concentrate.

The fluidized bed was operated at approximately 1000° C., and theconcentrate was fed at about 140 kg/h. Gas velocities of about 0.4 m/swere used. The residence time was rather low, at approximately 20seconds per pass. Most of the material underwent two passes through thereactor, with a small quantity passing through three times. The sulphurlevel decreased from 4.55% S to 0.5% after the first pass (96%elimination of S), and to 0.24% after the second pass (98% elimination),and to 0.13% S after the third pass. During roasting, the impuritieswere diminished as follows:

S from 4.55% to 0.24% (to 0.13%)

As from 40 to 21 ppm

Se from 60 to 8.8 ppm

Te from 10 to 7.8 ppm

Os from 5.5 to 3.8 g/t

Smelting was carried out in a pilot-scale DC arc furnace. 24 tons of(mostly double pass) dead-roasted concentrate (including 1 ton oftriple-roasted material) was processed in a week-long campaign. Thefurnace was operated at a power level of 300 to 500 kW, which translatesto a power flux of 290 to 480 kW/m². The average operating temperaturewas 1650° C. Calcine was fed to the furnace at feedrates of 200 to 300kg/h, and approximately 5% coke addition was used. No additional fluxeswere added. An energy requirement of 650 kWh/t of calcine was required(neglecting energy losses from the furnace shell). (Obviously in afull-scale plant operating with hot feeding of calcine to the furnace,this figure would be less.) The process was operated consistently withless than 1 g/t PGM in slag, and values as low as 0.3 g/t in the slagwere demonstrated. The average PGM loss to the slag over the entirecampaign was 2.9 g/t.

The analyses of the original concentrate, roasted concentrate, and slagare shown below (mass %).

Al₂O₃ C CaO Co Cr₂O₃ Cu Original concentrate 4.2 — 4.4 0.06 2.6 1.04Roasted concentrate 5.4 0.09 4.3 0.06 2.7 1.01 Slag 7.1 0.03 5.2 0.072.8 0.13 PGM, FeO MgO Ni S SiO₂ g/t Original concentrate 16.4 20.4 1.914.55 42.6 308 Roasted concentrate 16.3 19.9 1.84 0.25 43.3 296 Slag  7.224.7 0.10 0.07 51.0 2.9

Impurity removal overall (including roasting and smelting) is shownbelow, as a percentage of the amount originally present in the unroastedconcentrate.

Impurity removal in roasting and smelting, % of element in feed

As Bi Mn Pb Se Te V 70 87 95 100 95 84 77

Approximately 109 kg of alloy per ton of roasted concentrate wasproduced in the furnace. Over the campaign, about 2.6 tons of alloy wasproduced in total. Most of the alloy was tapped in two large batches.(The first alloy tap was diluted somewhat by the initial metal heel inthe furnace.) Shown below is the composition of the alloy, together withthe composition of the alloy produced in a laboratory-scale preliminarytest (all in mass %). Also shown is the composition of the refined alloyproduced by blowing the molten alloy with air, as discussed below.

C Co Cr Cu Fe Ni S Si PGM Small-scale 0.05 0.55 0.27 9.81 70.6 17.1 2.00<0.05 0.2804 test 836 kg alloy 1.06 0.33 3.35 7.56 67.7 16.6 0.48 1.340.1700 1612 kg alloy 0.97 0.50 2.35 7.43 71.1 15.3 0.97 1.05 0.2646Refined alloy 0.04 0.6 0.03 13 60 24 0.4 <0.05 0.2609

The alloys produced during the furnace campaign had the following rangesof composition.

C: 0.6-1.1%

Cr: 1.6-3.35%

Si: 0.76-1.34%

The alloy with the worst composition (i.e. from the 836 kg batch) wasselected to demonstrate the downstream process on the most conservativebasis. In order to lower the quantities of carbon and silicon (andchromium) prior to leaching, it was necessary to blow air into themolten alloy (using a top-blown rotary converter, to simulate theoperation of the proposed ladle furnace to be used for this operation).The composition of the resulting refined alloy is shown in the tableabove. This alloy was water-atomized to a particle size less than 100μm. The atomized alloy was then used for the leaching tests.

After hydrometallurgical processing, a final PGM concentrate wasproduced with the composition below (mass %).

PGM + Au Pt Pd Rh Ru Ir Au 61.4 34.6 12.7 4.3 7.7 1.67 0.61 Fe Ni Cu SiCr Se Te S C As 3.6 0.27 3.3 1.75 2.5 0.016 0.007 2.0 0.92 0.49

MatteRoast

Small-scale laboratory tests were carried out on PGM-containing furnacematte. The matte was either milled as a solid, or water-atomized fromthe liquid state, then dead-roasted in either a fluidized bed or arotary kiln. (No difference was found between the roasting behaviour ofthe milled and the atomized matte.) It was shown that matte can beroasted to extremely low levels of sulphur. The dead-roasted matte wasthen smelted in a two-stage process. The first-stage of smeltingproduces a small quantity of copper-nickel alloy that contains almost noiron or sulphur. The PGMs essentially all report to the copper-nickelalloy. The first-stage alloy has a PGM content around 2%. (This can beupgraded by leaching the Cu and Ni to produce a PGM concentrate.) Theslag from the first stage is then smelted, using a carbonaceousreductant, to produce a second alloy containing most of the remainingbase metals, as well as the residual precious metals.

Small-scale fluidized-bed roasting tests were carried out on 20 gsamples in a 25 mm silica tube fluidized bed. Successful roasting wasachieved using a particle size range of 250-300 μm, and a temperature of800-850° C. The sulphur content of the matte decreased from 28.7% to0.03% in 3 hours. Good results were also obtained after 1.5 hours at950° C. Temperatures above 900° C. are recommended for completedesulphurization. In order to provide larger samples for smelting testsfurther roasting was carried out in a laboratory-scale rotary kiln. A 20kg sample of furnace matte was crushed to a 100-600 μm particle sizerange. The roasting was accomplished in 4 days of operation, with 9passes of 12 hours each, with a stepped increase in temperature from675° C. to 1000° C. over this period. The sulphur content of thismaterial decreased from 26.7% to 0.04% by mass. A mass balance showsthat 1. kg of dead-roasted matte is produced from 1.05 kg of furnacematte as originally supplied. During roasting, the impurities werediminished as follows: S from 27.4% to 0.035%; Se from 244 ppm to 14ppm; Te from 96 ppm to 32 ppm; and As from 54 ppm to 46 ppm. There is nosignificant loss of PGMs, except for some Os.

A crucible test in a laboratory-scale furnace was performed using a feedcomprising 1050 g of dead-roasted furnace matte (derived from 1098 g ofunroasted furnace matte), 450 g of silica, and 31.5 g of carbon. Thisproduced 1517 g of slag, and 38 g of a copper-nickel alloy containingthe vast majority of the precious metals. The metal button that wasproduced was equivalent in mass to 12% of the Cu—Ni content of theoriginal furnace matte. This alloy quantity is comparable to the amountof PGM-containing alloy produced in the traditional slow-coolingprocess. The recovery of the precious metals was 99.0%, expressed as(PGM+Au in alloy)/(PGM+Au in alloy and slag).

PGM, Cu Ni Co S Fe FeO SiO₂ g/t Furnace 10.2 17.6 0.66 27.4 39.8 — — 762matte Roasted 10.7 18.1 0.68 0.035 41.9 — — 801 matte First-stage 65.331.l 0.048 0.37 <0.2 — — 19815 alloy First-stage 5.55 11.7 0.42 0.006 —37.3 29.2 5.2 slag

It is clearly quite possible to treat the slag from the first smeltingstage according to standard slag-cleaning practice in a DC arc furnace.Very high recoveries of the base metals and the residual precious metalswould be expected in the second-stage collection.

Zinc

Calcined zinc concentrate was fed, together with coke as a reductant, toa pilot-scale DC arc furnace, fuming off zinc vapour. (Other works, hasdemonstrated the production of Prime Westem grade zinc by furthertreatment of the zinc vapour in a lead splash condenser. It is alsopossible to use distillation to refine this zinc even further.)

A total of 56 tons of calcine was processed during the test work, withcoke and lime additions averaging approximately 13% and 3%,respectively. Approximately 16 tons of discard slag and 38 tons of zincoxide-rich bag-plant dust (fume) was produced by operating the DC arcfurnace at a power level between 500 and 700 kW. In this series oftests, the zinc vapour leaving the furnace was combusted with air andcollected in a bag plant. The feed materials included unagglomeratedcalcine, pellets dried to 150° C., pellets dried to 350° C., and pelletsindurated at 1300° C. The sulphur content of the feed materials variedbetween 1 and 2.4%.

The addition of between 12 and 13% coke resulted in an overall zincextraction efficiency of 95.4%. Fuming rates of up to 170 kg Zn/h per m²of bath area were obtained. (In the case of feeding unagglomeratedcalcine, a zinc extraction efficiency of 98.7% was obtained, and theaverage zinc fuming rate was 164 kg/h·m².)

The specific energy requirement was found to be approximately 1.17MWh/ton feed at an average operating temperature of 1490° C. Ironproduction varied between 2 and 21 kg per ton of calcine.

The fume produced during the test work was of an even better qualitythan that for previous test work during which the condenser wassuccessfully coupled to the furnace. The ratio of CaO, MgO, SiO₂, andFeO to ZnO was found to be approximately 0.04 in this test work,compared to a value of 0.14 found previously. Therefore it is reasonableto expect good condenser performance.

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What is claimed is:
 1. A process for treating a metal sulfide concentrate which includes at least one metal selected from the group consisting of the PGMs, nickel, cobalt and zinc, the method comprising the steps of: dead-roasting the metal sulfide concentrate, smelting the dead-roasted concentrate under reducing conditions in an electrically stabilized open-arc furnace, and collecting any metals from the smelting step in the form of an alloy or vapor.
 2. The process of claim 1 further comprising the step of: implementing the dead-roasting step to produce a steady stream of SO₂-bearing gas.
 3. The process of claim 2 further comprising the step of: using the SO₂-bearing gas from the dead-roasting step as a feedstock in a sulfuric acid plant.
 4. The process of claim 3 further comprising the step of: scrubbing and neutralizing the SO₂-bearing gas produced by the dead-roasting step.
 5. The process of any of claims 2-4 further comprising the step of: performing the dead-roasting step in an enclosed vessel to avoid unwanted dilution by air of SO₂.
 6. The process of claim 5 further comprising the step of: conducting the dead-roasting step in a fluidized bed reactor.
 7. The process of claim 1 further comprising the step of: introducing coke in the smelting step to achieve said reducing conditions.
 8. The process of claim 1 wherein the concentrate includes zinc and which further comprises the steps of: releasing zinc from the metal sulfide concentrate during the dead-roasting step, melting the released zinc under the reducing conditions of the smelting step to form a zinc metal vapor, fuming off the zinc metal vapor in a gas stream, and recovering the zinc metal by condensation of the gas stream.
 9. The process of claim 8 further comprising the step of: agglomerating the zinc released during the releasing step to form a concentrate before smelting the dead-roasted concentrate.
 10. The process of claim 1 further comprising the step of: using a PGM concentrate for said metal sulfide concentrate.
 11. The process of claim 10 further comprising the step of: using a green furnace matte for said PGM concentrate.
 12. The process of claim 1 wherein the concentrate includes iron and which further comprises the step of: removing iron in oxide form from said alloy.
 13. The process of claim 1 wherein the feed to the furnace includes at least one member selected from the group consisting of carbon, silicon and chromium, and which further comprises the step of: removing said at least one member selected from the group consisting of carbon, silicon and chromium from said alloy.
 14. The process of claim 13 further comprising the step of: performing the removing step in a converter.
 15. The process of claim 14 further comprising the step of: atomizing alloy from the converter in a form that is suitable for subsequent hydrometallurgical recovery of metal.
 16. The process of claim 1 wherein the smelting step comprises: a two-stage reduction smelting process.
 17. The process of claim 16 further comprising the steps of: operating a furnace under slightly reducing conditions in a first stage, and operating said electrically stabilized open arc furnace under highly reducing conditions in a second stage.
 18. The process of claim 17 wherein the concentrate includes iron and which further comprises the steps of: producing in the second stage an iron-based alloy, atomizing the iron-based alloy, and subjecting the atomized alloy to hydrometallurgical treatment.
 19. The process of any of claims 16-18 wherein the concentrate contains copper and nickel and wherein the first stage comprises the steps of: producing a copper/nickel alloy, forming particles of the copper/nickel alloy by at least one of water atomization, granulation, or crushing and milling, and subjecting the particles of copper/nickel alloy to hydrometallurgical treatment. 